Process for smelting copper from mixed copper-zinc-lead materials



United States Patent 3,231,376 PRGCESS FOR SMELTENG CGPPER FRQM MIXED tCtBPPER-ZiNC-LEAD MATERIALS Stephen lhiiliam Kenneth Morgan, Bristol, England, as-

signor to Metallurgical Processes Limited, Nassau, Bahamas, a corporation of Bahamas, and The National smelting ompany Limited, London, England, a British company Filed Sept. 19, 1962, er. No. 224,762 Gaines priority, application Great Britain, Sept. 25, 1961, 34,263/61 9 Claims. (Cl. 75-74) This invention relates to the separation of copper from oxidic materials containing zinc, lead and copper by treatment in a zinc blast furnace.

The primary purpose of this process is the recovery in saleable form of zinc metal and lead metal but it is known that copper can also be obtained in a form that is suitable for further treatment by well-known methods to produce copper metal.

The main component of the charge of a zinc blast furnace other than the carbonaceous fuel is usually an oxidic lump material such as is obtained for instance by sinterroasting on a Dwight Lloyd machine, which contains the bulk of the zinc to be smelted. As the zinc blast furnace and particularly the condenser unit and lead circuits generate considerable quantities of dusts and drosses rich in zinc and lead it is convenient to recover these metals by incorporating such materials in the furnace feed, for instance, by mixing with the charge to the Dwight Lloyd machine. Thus the charge to the furnace preferably always contains some lead, most of which is recovered as metal: mixed lead and zinc concentrates may thus be advantageously smelted. This lead, which is tapped together with the slag from the furnace, has the desirable property of collecting any small amounts of silver, gold and copper present in the charge. In the absence of any lead fall to the furnace bottom these valuable metals would almost certainly be lost.

It is known that substantial recoveries of copper incorporated in the sintered zinc-containing lump material (hereinafter referred to as sinter) in the form of a matte can be achieved when the weights of lead and copper charged are nearly equal. Small amounts of copper are also contained in the lead.

Even if only a small amount of copper is present, not all of it is tapped with the lead; the copper always distributes itself between the lead phase and a matte phase. This is understandable as, apart from any sulphur contained in the sinter, commonly used carbonaceous fuels contain some sulphur.

Thus, Welsh metallurgical cokes contain about 1% sulphur and this would be enough with the l2% copper in the sinter to form cuprous sulphide. On the other hand, it is known that copper and lead are miscible in all proportions at the temperature of the hearth of the furnace, that is, above the critical point (990 C.) of this alloy system, so that all the copper might have been expected to dissolve in the lead. There is not enough information available on the system Pb-Cu-S to enable predications on the course of events in the furnace to be made.

The above discussion is relevant to the substance of the present invention. For whereas metals (lead and copper) and slag difier in density by about 7 gm./millilitre, so that they readily separate into two layers or may be readily separated, the difference in density between a matte and slag is much lessabout 1 gm./ml. Thus any effect which tends to hinder the settling of the denser liquid phase will affect the settling of matte more than that of metal. It is well known that in copper smelting the copper content of slags is greater than the equilibrium value. According to copper smelters these high copper 3,231,379 Patented Jan. 25, lgfifi contents of reverberatory slags are attributed to suspended matte and this lack of settling is in part, at least, due to the presence of crystals of magnetite in the slag.

I has now been discovered that in the zinc blast furnace smelting of mixed copper-lead-zinc oxidic materials in which part of the copper is recovered as a matte, losses of copper in slag greater than that due to the solubility of copper or its compounds are in part brought about by crystals of various sorts in the slag. It has further been discovered that these copper losses can be reduced by dissolving these crystals by the addition of small amounts of fluxes, especially silica, in a manner to be described. Furthermore, these additions do not affect the recovery of zinc unfavourably.

At all points in the shaft of the furnace in the vicinity of the tuyeres, the desirable and necessary reactions occurring between the gas and the particles or lumps of sinter, namely, the net addition of heat to the sinter and loss of zinc therefrom, will proceed faster if each lump of sinter is wholly solid than when each is covered with a film of slag, however thin. When this unfavourable condition is reached, the reactions still continue, though more slowly, so that eventually so much of each separate piece of sinter is molten that this slag, which is molten sinter gangue plus some zinc oxide, flows over the coke or other carbonaceous fuel and falls below the tuyere zone. There, no further reaction takes place as below the tuyeres there is no gas flow.

Thus, the longer the onset of melting is delayed by increasing the solidus (melting point) temperature of the sinter gangue, the hotter the slag eventually becomes and the greater the net volatilisation of zinc. The slag that is finally tapped consists of the sinter gangue plus some dissolved zinc oxide plus the carbonaceous fuel ash and any flux that was added as part of the furnace charge and which is soluble in the slag. Obviously, any flux or fuel ash cannot aflfect the extent of zinc recovery since they cannot dissolve in the usually zinciferous sinter gangue slag until the latter has begun to melt, unless these additives have a melting point below that of the sinter gangue.

T he freezing point or liquidus of the slag that is tapped, hereinafter called the final slag, is not important. All that is required of the final slag is that it should run out of the furnace sufliciently freely. For this, a temperature of tapping above the freezing point or liquidus is desirable but not essential: the presence of, say, 10% suspended crystals in the final slag need not prevent it from running freely out of the furnace.

However, the situation is quite otherwise if the slag is required to run out of the furnace into a forehearth or settler where lead and matte settle out and are separately collected. Firstly, the slag must stay fluid much longer, that is to say, it must be able to stand a drop in temperature without immediately freezing. Secondly, the slag must not be mushy or contain suspended solids which prevent small droplets of matte settling out or coagulating to form bigger droplets. Therefore, the slag must be fully molten and preferably be at a temperature of 50-109" C. above its freezing point or liquidus. At the same time the final slag and other compositions must be so chosen as to give good zinc recoveries as outlined in the previous three paragraphs; having as they must high solidus temperatures, these in general have high liquidus temperatures also.

The invention consists in a method of treating the slag run from a zinc blast furnace smelting copper-lead-zinc oxidic materials which consists in adding a non-alkaline (acid or neutral) flux in lump form to the furnace burden, the flux having itself a melting point high enough for it to pass through the furnace substantially unrnclted until it commingles with the sinter gangue in the slag pool below tuyere'level, where it lowers the liquidus temperature of the slag in order to facilitate separation of matte and slag, e.g. in a forehearth.

Such a flux addition does not harm the zinc recoveryas it would if it were made to the sinter mix, since fluxes which lower the liquidus normally also lower the solidus provided the flux has a high melting point. The effect of such additions upon the liquidus of the mixture slag plus flux depends on the composition of the slag and the amount of flux used. Suitable fluxes are silica of alumina in lump form, or iron oxide or iron, or mixtures of any or all of these.

The form in which iron is used is preferably scrap metal, since iron will assist the reduction of zinc oxide in the slag; hematite or magnetite would increase the carbon dioxide content of the furnace: such an increase might affect the recovery of zinc.

Alkali-containing fluxes, or fluxes melting below 1,100 0., are not suitable. The acid or neutral flux must be in lump form and not powders so that it passes relatively rapidly through the furnace to commingle with the sinter gangue substantially only in the slag pool below the level of the tuyeres.

Preferably for the same reason, the fiux should be added separately from the sinter, for example, through a separate device.

During the operation of the furnace, the direction of flow of the furnace burden is generally vertically downwards, with little or no sideways movement of the charge. When the flux is charged at the top of the furnace, if it is charged to one side of the main mass of charge (sinter mix, coke, etc.). there will be very little tendency for the lumps to mix with the rest of the charge through lateral movement. According to a preferred form of our invention, therefore, the lump flux may be charged to the furnace at the top through a separate charging device sit- For the purposes of the calculation, we assume a constant 37% Zn in sinter, an amount of coke in the charge such that the ratio Zn/C (in fuel) is 1.30 and that the composition of the coke ash is (as percent of carbon in coke) 4.10% SiO 2.90% A1 1.00% Ref), 0.40% CaO.

These calculations for a random set of sinter and slag compositions illustrate that additions of silica and in a few cases, silica plus iron, of between five and ten percent of the slag weight reduce the slag freezing point to a temperature about the same as that at which the initial sinter gangue begins to melt. Thus the slags from which the copper-containing mattes, and lead, have to segregate are fully molten and, in fact, possibly 100 C. above the slag freezing point.

Other calculations on the compositions cited above show that other additions, for instance iron oxide or alumina, either have no marked effect or raise the liquidus temperature. Nevertheless there are slag compositions where additions of iron oxide or alumina would lower the liquidus of the slag. Such slags might arise, for instance, where the fuel ash has a composition and volume much different from that used in the example.

Results of the calculations involved are shown in the table below.

Various modifications may be made within the scope of the invention.

The accompanying drawing shows a section through a zinc blast furnace having a slag tapping point 2 at the bottom and a gas (zinc vapour) off-take 3. The furnace is charged by means of two principal bell charging devices 4 and the flux according to the present invention is charged by means of a separate smaller bell charging device 5 in an amount corresponding to 1% to 2% of the sinter weight. The tuyere level of the zinc blast furnace is indicated by the broken dash lines 6, the slag pool by the' numeral 7, and the forehearth or settler by the numeral 8.

Composition of Sinter Sinter Gangue lilnal Flux Added Flux Added 8 Example Liquidus Total (N 0 Percent Per- Percent Per- Iron 6210 S A1203 MgO Solidus Liquidus Fluxes of cent Liquidus of cent Liquidus as FeO Added) Sinter Slag Sinter Slag 13. 1 8. 0 4. 3 7. 3 0 1, 275 1, 400 1, 350 1% S10: 2. 8 1, 310 2% S102 5. 7 1, 300 3. 2 8. 6 7. 5 3.0 0 1, 300 1, 340 1, 280 {1% S10: 8. 1 1, 250 2% S10; 16. 2 1, 200 1% FeO 2% FeO 5. 8 7. 3 5. 5 0 0 1, 223 1, 500 1, 210 1% $10 1 9.5 1,180 2% S10 l9. 0 1, 170

1% F00 j 2% F eO Q. 8 9. 3 6. 5 1. 6 0 1, 160 1, 350 1, 230 1% S102 6. 8 1, 230 {2% S10 12. 6 1, 200

1% F60 2% E 8.9 6. 6 4. 5 0. 6 0 1, 155 1, 400 1, 210 1% S10; 4. 5 1,170 2% S10: 9. 1 1, 160 12. 2 10. 3 8.0 2. 6 0 1, 210 1, 270 1, 270 1% SiOg 2. 8 1,250 2% $102 5. 6 1, 210 5. 7 6. 4 4. 5 2. 2 0 1, 240 1, 300 1, 270 1% SiOg 4. 7 1, 240 2% S102 9. 5 1,160 12. 1 5. 3 5. 0 1. 1 O 1, 180 1, 245 1, 230 1% S10; 3. 9' 1,170 2% S10; 7. 8 1, 160 9. 0 7. 5 6. 0 1. 2 0 1, 150 1, 270 1, 210 1% S10; 3. 8 1, 170 2% S102 7. 7 1, 120 9. 8 7. 3 5. 6 2. 7 2. 4 1, 210 1, 350 1, 420 3% $101 9. 9 1, 270

uated to one side of the furnace, preferably directly above the tapping for the slag. Thus the flux may travel substantially the vertical length of the furnace, without intermingling with the rest of the charge and, therefore, without consequent detrimental effects on the furnace behaviour, but having been heated to a sufiiciently high tem perature to facilitate its dissolution in the molten sinter gangue. A further factor which serves to separate flux and charge inside the furnace is that the flux is charged only intermittently and at different times from the main charging.

The invention will be further described with reference to the following example:

EXAMPLE To illustrate the invention, we consider sinters containing -40% Zn, 10% Pb, 1l0% Cu and varying amounts of gangue constituents. These are smelted with coke, the composition of the final slag thus calculated and the liquidus temperature deduced. The same calculations are then made for the case in which fluxes are added to the furnace burden, either 1 or 2% of the sinter weight.

What I claim is:

1. In the zinc blast furnace smelting of a charge containing oxides of zinc, lead and copper, together with carbonaceous reducing fuel, in which the zinc oxide is reduced and volatilized and the resulting vapor is withdrawn from the upper portion of the furnace shaft; the lead and copper oxides are reduced in large part and collected in a molten pool of slag, molten lead and copper matte at the bottom of the shaft; the slag is tapped from the bottom of the shaft into a body of molten slag; and molten lead and copper matte are settled out and separately recovered from the slag, the improvement in combination therewith which comprises:

(a) moving sintered lumps of such a charge of zinc, lead and copper oxides in the form of a column by gravity, downwardly, through the furnace shaft; the charge, however, containing a valuable and important amount of copper;

(b) smelting simultaneously the charge lumps of zinc, lead and copper oxides during their descent in the furnace shaft;

(c) adding separately to the upper portion of the charge column regulated amounts of a flux, also in lump form, selected from the group: silica, alumina, iron oxide and iron;

(d) using flux lumps having a melting point high enough to pass through the furnace shaft unmelted;

(e) advancing the flux lumps in the form generally of an inner column surrounded at least for the most part by the outer column of sintered charge lumps downwardly through the furnace shaft;

(f) passing the inner column of flux lumps downwardly in the shaft in unmelted form while surrounded by the outer column of charge lumps and while the lumps of the outer column undergo smelting to reduce in large part the oxides of zinc, lead and copper;

(g) moving simultaneously the inner column of flux lumps and the outer column of charge lumps sideby-side with little lateral movement downwardly in the shaft by gravity with little mixing of the flux lumps with the charge lumps;

(h) dropping the resulting hot gangue from the smelted outer column of charge lumps into the molten pool collecting at the foot of the furnace shaft;

(i) preheating the inner column of flux lumps during its passage downwardly in the shaft (1) to a temperature sufficiently low to inhibit reaction between the flux lumps and the surrounding charge lumps while only the charge lumps undergo smelting above the tuyere level of the furnace, but (2) to a temperature sufliciently high to facilitate dissolution of the flux lumps in the molten sinter gangue collected in the molten pool at the bottom of the furnace;

(j) melting the unmelted flux lumps by dropping and commingling them with the hot sinter gangue collecting in the molten slag pool below the tuyere level of the shaft;

(k) adding enough of the preheated flux lumps in the pool to lower the freezing point, and hence increase the fluidity, of the molten slag to facilitate cleaner and sharper separation of the slag from copper matte;

(1) tapping from the pool the resulting high fluidity molten slag containing the molten lead and copper matte; and

(m) settling the molten mixture so withdrawn to separate the molten lead and copper matte from the slag.

2. Method according to claim 1, adding the flux lumps to the column of charge lumps directly above the place from which the slag pool is tapped to facilitate passage of the inner column of flux lumps through the furnace shaft to the slag pool and to inhibit mixing of the flux lumps with the charge lumps.

3. Method according to claim 11, passing the inner column of flux lumps rapidly through the furnace shaft to commingle with the sinter gangue in the slag pool below the level of the tuyeres.

4. Method according to claim 1, adding the flux lumps to the charge intermittently to inhibit more effectively mixing of the inner column of flux lumps with the outer column of charge lumps inside the furnace shaft.

5. Method according to claim 1, adding the flux lumps to the charge lumps at the side of the column of charge lumps in the furnace shaft.

6. Method according to claim 1, in which the melting point of the flux lumps is above 1100 C.

7. Method according to claim 1, in which the copper content of the sintered charge lumps is 110% by weight.

8. Method according to claim 1, delaying the onset of melting of the sintered charge lumps by increasing the solidus temperature of sinter gangue present in the charge to increase the temperature of the slag and the volatilization of zinc.

9. Method according to claim 1, in which the slag tapped from the slag pool is fully molten and at a temperature sufiiciently above its liquidus to be substantially free of suspended solids to inhibit small droplets of matte from settling out to form larger droplets.

References Cited by the Examiner UNITED STATES PATENTS 805,835 11/1905 Baggaley -74 2,598,743 6/1952 Waring 7586 2,693,410 11/1954 Waring 7586 2,802,661 8/1957 McCutcheon 26627 2,932,566 4/1960 Lumsden 7592 3,073,696 1/1963 Lumsden et al. 7587 3,139,472 6/1964 Evans 26627 DAVID L, RECK, Primary Examiner. 

1. IN THE SIZE BLAST FURNACE SMELTING OF A CHARGE CONTAINING OXIDES OF ZINC, LEAD AND COPPER, TOGETHER WITH CARBONACEOUS REDUCING FUEL, IN WHICH THE ZINC OXIDE IS REDUCED AND VOLATILIZED AND THE RESULTING VAPOR IS WITHDRAWN FROM THE UPPER PORTION OF THE SURFACE SHAFT; THE LEAD AND COPPER OXIDES ARE REDUCED IN LARGE PART AND COLLECTED IN A MOLTEN POOL OF SLAG, MOLTEN LEAD AND COPPER MATTE AT THE BOTTOM OF THE SHAFT; THE SLAG IS TAPPED FROM THE BOTTOM OF THE SHAFT INTO A BODY OF MOLTEN SLAG; AND MOLTEN LEAD AND COPPER MATTE ARE SETTLED OUT AND SEPARATELY RECOVERED FROM THE SLAG, THE IMPROVEMENT IN COMBINATION THEREWITH WHICH COMPRISES: (A) MOVING SINTERED LUMPS OF SUCH A CHARGE OF ZINC, LEAD AND COPPER OXIDES IN THE FORM OF A COLUMN BY GRAVITY, DOWNWARDLY, THROUGH THE FURNACE SHAFT; THE CHARGE, HOWEVER, CONTAINING A VALUABLE AND IMPORTANT AMOUNT OF COPPER; (B) SMELTING SIMULTANEOUSLY THE CHARGE LUMPS OF ZINC, LEAD AND COPPER OXIDES DURING THEIR DESCENT IN THE FURNACE SHAFT; (C) ADDING SEPARATELY TO THE UPPER PORTION OF THE CHARGE COLUMN REGULATED AMOUNTS OF A FLUX, ALSO IN LUMP FORM, SELECTED FROM THE GROUP: SILICA, ALUMINA, IRON OXIDE AND IRON; (D) USING FLUX LUMPS HAVING A MELTING POINT HIGH ENOUGH TO PASS THROUGH THE FURNACE SHAFT UNMELTED; (E) ADVANCING THE FLUX LUMPS IN THE FORM GENERALLY OF AN INNER COLUMN SURROUNDED AT LEAST FOR THE MOST PART BY THE OUTER COLUMN OF SINTERED CHARGE LUMPS DOWNWARDLY THROUGH THE FURNACE SHAFT; (F) PASSING THE INNER COLUMN OF FLUX LUMPS DOWNWARDLY IN THE SHAFT IN UNMELTED FORM WHILE SURROUNDED BY THE OUTER COLUMN OF CHARGE LUMPS AND WHILE THE LUMPS OF THE OUTER COLUMN UNDERGO SMELTING TO REDUCE IN LARGE PART THE OXIDES OF ZINC, LEAD AND COPPER; (G) MOVING SIMULTANEOUSLY THE INNER COLUMN OF FLUX LUMPS AND THE OUTER COLUMN OF CHARGE LUMPS SIDEBY-SIDE WITH LITTLE LATERAL MOVEMENT DOWNWARDLY IN THE SHAFT BY GRAVITY WITH LITTLE MIXING OF THE FLUX LUMPS WITH THE CHARGE LUMPS; (H) DROPPING THE RESULTING HOT GANGUE FROM THE SMELTED OUTER COLUMN OF CHARGE LUMPS INTO THE MOLTEN POOL COLLECTING AT THE FOOT OF THE FURNACE SHAFT; (I) PREHEATING THE INNER COLUMN OF FLUX LUMPS DURING ITS PASSAGE DOWNWARDLY IN THE SHAFT (1) TO A TEMPERATURE SUFFICIENTLY LOW TO INHIBIT REACTION BETWEEN THE FLUX LUMPS AND THE SURROUNDING CHARGE LUMPS WHILE ONLY THE CHARGE LUMPS UNDERGO SMELTING ABOVE THE TUYERE LEVEL OF THE FURNACE, BUT (2) TO A TEMPERATURE SUFFICIENTLY HIGH TO FACILITATE DISSOLUTION OF THE FLUX LUMPS IN THE MOLTEN SINTER GANGUE COLLECTED IN THE MOLTEN POOL AT THE BOTTOM OF THE FURNACE; (J) MELTING THE UNMELTED FLUX LUMPS BY DROPPING AND COMMINGLING THEM WITH THE HOT SINTER GANGUE COLLECTING IN THE MOLTEN SLAG POOL BELOW THE TUYERE LEVEL OF THE SHAFT; (K) ADDING ENOUGH OF THE PREHEATED FLUX LUMPS IN THE POOL TO LOWER THE FREEZING POINT, AND HENCE INCREASE THE FLUIDITY, OF THE MOLTEN SLAG TO FACILITATE CLEANER AND SHARPER SEPARATION OF THE SLAG FROM COPPER MATTE; (L) TAPPING FROM THE POOL THE RESULTING HIGH FLUIDITY MOLTEN SLAG CONTAINING THE MOLTEN LEAD AND COPPER MATTE; AND (M) SETTLING THE MOLTEN MIXTURE SO WITHDRAWN TO SEPARATE THE MOLTEN LEAD AND COPPER MATTE FROM THE SLAG. 